Direct production of refined metals and alloys

ABSTRACT

The present invention provides a method of reducing metal oxide material to metal comprising the steps of forcibly circulating molten carrier material in a closed loop path serially through a charge reduction zone on one arm of the loop, a combined melt desulphurisation zone and post combustion or heating zone on the other, reducing said metal oxide to solid metal by the carbonaceous material contained within a mixed composite charge of the metal oxide, carbonaceous reductant and flux in said reduction zone, the metal oxide and carbonaceous reductant being utilised in proportions such that the carbon from the carbonaceous reductant is converted to carbon monoxide; reacting carbon monoxide with oxygen in the refining loops downstream from the reduction loop before being combusted to completion in the heating zone at the surface of the molten carrier material so that heat generated by the reaction is transferred to the molten carrier material which is circulated to the reduction zone; separating a metallised raft from said molten carrier material by projecting said metallised raft along into the first refining loop by virtue of the drag force exerted on the metallised raft by the circulating carrier material before the molten carrier material is circulated to the heating zone so that the surface of the molten carrier material which is circulated to the heating zone is substantially free of solid material.

This invention relates to thermal reduction of metal oxide materials,for example iron oxide ores such as haematite, metal oxide ores, e.g.nickel-laterite and chromite ores, and deepsea manganese nodules, forthe purpose of recovering metallic values therefrom in a refinedcondition on a continuous basis. The present invention is also concernedwith the treatment of steel plant fines, in-plant fines and othersecondary materials generated during metallurgical and wasteincineration operations.

A method for recovery of metals from oxide materials is disclosed in EP0266975 and U.S. Pat. No. 4,701,217, in which smelting reduction ofmetallic oxide materials, e.g. pelletised iron oxide ore or iron orefines is effected by contacting such material with a circulating moltencarrier material, e.g. molten iron in the case of iron oxide smelting,in a furnace, and by introducing a carbonaceous reductant e.g. coal intothe carrier material. The reductant converts the metal oxide to metal,which is assimilated immediately into the molten carrier material in asmelting reduction zone. Carbon monoxide thereby produced is combustedin a heating zone through which the carrier material passes to effectheat recovery. Slag is removed from the carrier material before enteringthe heating zone. A protective layer of molten material, e.g. lead,which is substantially stationary is maintained below the circulatingcarrier material to inhibit erosion of the furnace hearth.

Further details of oxide ore smelting by the above melt circulationmethod are disclosed in “Overview of generic melt circulationtechnology” by N A Warner in Proceedings of the International Symposiumon Challenges in Process Intensification, Montreal, Quebec Aug. 24-29,1996, The Canadian Institute of Mining, Metallurgy and Petroleum;269-281, 1996; (ISBN 0-919086-68-3).

EP 0266975 and U.S. Pat. No. 4,701,217 also make reference to laboratorytestwork in which cylindrical compacts of mixed pulverised coal andhaematite ore fines were immersed in a liquid metal heat transfer medium(lead) for two minutes. Below 900° C. very little reaction occurred,whereas at 1300° C. metallisations as measured by bromine-methanoldissolution were in the range 85-90%. It was suggested that if thecomposite pellets were immersed in hot metal (molten iron containingdissolved carbon) even faster rates could be expected. Theseobservations were not actually made use of in the processes discussed inthese former patents but now they are germane to the present invention.

It has now been discovered that the use of a circulating molten carriercan be advantageous not only in an improved method for reduction ofoxide of certain metal oxides but also to recover on a continuous basisrefined metal or alloy therefrom in preparation for continuous castingor other downstream finishing operations, e.g. ingot casting,fragmentation, globulation, granulation or if a powered product isrequired, atomisation.

According to the present invention, there is provided a method ofreducing metal oxide material to metal comprising the steps of forciblycirculating molten carrier material in a closed loop path seriallythrough a charge reduction zone on one arm of the loop, a combined meltdesulphurisation zone and post combustion or heating zone on the other;reducing said metal oxide to solid metal by the carbonaceous materialcontained within a mixed composite charge of the metal oxide,carbonaceous reductant and flux in said reduction zone, the metal oxideand carbonaceous reductant being utilised in proportions such that thecarbon from the carbonaceous reductant is converted to carbon monoxide;reacting the carbon monoxide with oxygen in the refining loopsdownstream from the reduction loop before being combusted to completionin the heating zone at the surface of the molten carrier material sothat heat generated by the reaction is transferred to the molten carriermaterial which is circulated to the reduction zone; separating ametallised raft from said molten carrier material by projecting saidmetallised raft along into the first refining loop by virtue of the dragforce exerted on the metallised raft by the circulating carrier materialbefore the molten carrier material is circulated to the heating zone sothat the surface of the molten carrier material which is circulated tothe heating zone is substantially free of solid material.

In contra-distinction to the method described in EP 0266975 and U.S.Pat. No. 4,701,217 the molten carrier material is not involvedchemically to a major extent in the reduction process but is there totransport the charge and reaction product materials in the solid stateat elevated temperature through an extended path in contact with themolten carrier material which provides the endothermic heat requirementsfor chemical reactions to tale place within a layer of composite chargematerial, deposited across the width of the molten carrier material nearto one end of the charge arm. A chemical reaction front moves throughthe layer commencing with solid material submerged in the molten carriermaterial on the underneath side of the deposited layer, possiblyinvolving an initial formation of solidified crust of the molten carriermaterial when the charge materials are first deposited at the feed endand then as heat is conducted into the layer, the reaction frontprogressively moves upward through the layer as it floats away along thearm on the molten carrier material, all the time releasing coalvolatiles and gaseous products of reduction into the gas space above thelayer which is enclosed by walls and a roof of thermal insulatingmaterial. Except for minor addition of oxygen to control carbondeposition and to facilitate a minor degree of sintering of thecomposite charge surface, no major combustion takes place in this gasspace and the conditions throughout are reducing to the newly formedmetallic phase, so there can be no question of reoxidation of metallicproduct taking place in this arm, even if some areas are fully reduced,whilst others are still evolving reducing gases. Accordingly, it ispossible to achieve high levels of metallisation of the charge materialwithout elaborate charge preparation or having to reduce the intensityas metallisation approaches completion. The partial sintering of thecharge surface, referred to above, occurs at temperatures around 950° C.in advance of the somewhat higher temperatures associated with reductionreaching the top surface.

Conduction within agglomerated materials is the principal heat transfermeans even at very high temperatures and in the present case this isenhanced by metal not originally in the charge materials infiltratinginto the charge layer as production progresses to enhance that producedby reduction. Also sintering phenomena which result ultimately in theformation of a metallised raft emerging from the far end of the armcause a progressive increase in thermal conductivity the longer thematerial stays floating down the metallisation arm of the reductionloop. Direct contact with liquid metal eliminates the need for radiativeheat transfer in the metallisation arm so there is no specialrequirement to provide geometric arrangements that ensure access todirect thermal radiation as this is no longer an issue as it is withcurrent technology.

The capillary rise effect occurs as a consequence of hot metal wettingthe newly-formed solid metal as a result of gas phase reactions andbeing drawn into the porous structure by surface tension. From capillarytheory, the rise is directly proportional to the surface tension andinversely proportional to the density of the liquid, so this means thatfor molten iron the propensity for capillary rise is more than twice asgreat as that for water entering a porous solid material, assuming thatthe degree of wetting as measured by the so-called contact angle is thesame in both cases.

In the present invention, the infiltration of hot metal by the capillaryrise phenomenon as reduction proceeds, effectively reduces the thermalresistance between the molten carrier material and the progressivelyupwards moving reduction zone to such an extent that heat transferinfluences are very considerably reduced so that the kinetics of thechemical reactions become all important, a situation under the controlof the process designer in so far as the chemical kinetics can beregulated by specifying particle size of the reacting solids. This meansthat the most cost-effective solution can be identified for a particularcase.

In the absence of heat being transferred from above, as in currentpractice, the temperature of the top surface of the charge material asseen by the roof and wall enclosure at no time exceeds about 1300° C.,so the use of relatively inexpensive and low thermal mass insulatingmaterials becomes a practical proposition throughout the charge arm ofthe reduction loop. Furthermore, these materials will be unlikely to beexposed to the carry-over of semi-molten oxide components because thereaction gases are no longer passing through the highest temperaturezone in their passage to the gas space above the charge. Withestablished rotary hearth processing, for example, the situation isreversed. By all these means reduction rates achievable by the presentinvention are potentially faster than those associated with rotaryhearth furnaces, operated with more expensive refractories at highertemperature levels.

As a further consequence of no major combustion taking place in thecharge arm, the problem of fines being blown off the charge material andadhering to the wall of the furnace in a semi-molten condition causingtrouble to facilities, including accretion formation requiring periodicremoval, are very considerably reduced. The principal combustionprocesses all occur elsewhere in the overall circuit where finesgeneration is no longer a major issue. Also, the charge end of themetallisation arm is relatively cool in comparison with the temperaturesneeded from the outset in processes employing rotary hearth furnaces fortreating feed agglomerates, so if any solid accretions are formed in thefront end of the arm, these will not be partially fused and will beremovable with relative ease in any case. As the charge layer travelsdown the arm it will undergo progressive sintering and by the time it isadjacent to roof temperatures approaching the melting point of metallicoxides associated with the charge, fines generation from the sinteredsurface are most unlikely to occur. However, an open structure forrelease of coal volatiles and reaction product gases freely into the topgas space is maintained throughout.

As an additional measure to ensure that fines are not carried overexcessively into the gas flowing above the charge layer, the gasfreeboard distance above the solids may be arranged so that local gasvelocities are consistent with the propensity for solids entrainment.This means that in many cases the gas velocity has to be relatively lowin the charging region with progressive increase permissible as thesolids sinter or agglomerate partially, a condition achieved, forexample, by profiling the cross-section of the furnace gas space. In anyevent, even without such profiling it is desirable to arrange for thedischarge of evolved gases at the far end from the charge position sothat the increase in cumulative gas flow is in contact with solids withprogressively longer times to sinter or consolidate. All the abovefactors contribute to amelioration of potential problems with finesgeneration, one of the more serious shortcomings of current technology.

There is an additional possible scenario, which again flows principallyfrom the infiltration of molten metal into the charge layer as reactionproceeds. Not only is the thermal conductivity increased but also theaverage density of the contents of the floating raft is graduallyincreased as more and more metal is infiltrated by capillary action tofill up the voids beneath the principal reduction zone as it progressesupwards. This has the effect of causing the raft to sink further intothe molten carrier material and increases the amount of reduction thattakes place by true smelting reduction, involving the carbon dissolvedin the molten carrier material. In the limit, this behaviour means thatthe metallised raft contains considerably less metal than was originallypresent in the feed. In the extreme case, the metallised raft projectedout of the reduction loop is comprised of residual solid metal formed bygaseous reduction, which has not yet had the opportunity to beassimilated into the molten carrier material, in association with slagconstituents and excess coke coal all forming an agglomerated solidstructure. Under these circumstances, an increased amount of moltencarrier material clearly has to be overflown along with the metallisedraft to the steelmaking loop in order to balance any new metal beingadded to the molten carrier material in the reduction loop itself.

The mixed charge to the charge arm of the reducing loop is comprisedprincipally of carbonaceous material, metal oxide material and usually asuitable flux, e.g. lime, limestone or dolomite with particle sizerequirements determined by preliminary laboratory testing: In some casesit may be desirable to introduce a degree of compaction into the mixedcharge constituents as they are deposited onto the moving surface of themolten carrier material using means well-known to those skilled in theart. Also it may be advantageous in some cases to deposit a filter cakefrom a horizontal belt filter to be discharged onto the melt surfaceafter initial in-line drying. However, in many cases simple mechanicalmixing of the constituents will be all that is required.

The vigorous reaction and ebullition that might have taken place andcaused an unacceptable loss of fine material entrained in the evolvedgasses during smelting reduction of a very thin layer of depositedsolids is precluded to a large extent by depositing the composite chargesolids so that an immediate layer in the region of 5 to 10 cm thickeffectively physically blankets the surface of the molten carriermaterial, assisted by the increased thermal mass of the charge in somecases causing an initially frozen crust of molten carrier material onthe underneath side of the floating charge layer, slowing down vigorousreduction until the material at the interface reaches temperature of1100° C. or so by which time the charge solids have had an opportunityto consolidate at least to some extent.

The method of the present invention is usually performed with a verylarge proportion of molten carrier material circulation to metalproduced. For example, in the case where the metal oxide is iron oxide,a circulation ratio of 100:1 to 500:1 can be employed in the ironmakingloop, the actual ratio depending upon the nature of the feed and theenergy requirements thereof. In a particular example, a circulationratio of 315:1 is employed, namely for every unit of iron produced, 315units of molten iron as carrier are circulated between the reductionzone and the heating zone in the closed loop path. The rate ofcirculation depends upon the size of the equipment and the requiredreduction rate.

The metal oxide material incorporated within the composite mixed charge,which forms the layer floating on the molten carrier material, issubjected to reduction to produce the metal, and carbon monoxide isevolved above the surface of the molten carrier material and ispreferably passed to the other treatment stages which may optionallyinclude liquid metal based hot gas clean-up or other means ofdesulphurisation; boosting in pressure; staged partial combustion withoxygen to facilitate melting of the metallised raft and combustion ofexcess carbon therein; liquid slag formation; and decarburisation of themolten carrier material in at least one (preferably two) downstream meltcirculation loops by top blowing and direct flame impingementmechanisms, before being passed to the heating zone of the reductionloop where combustion to carbon dioxide is finally completed.

The temperature of the molten carrier material depends upon the type ofmetal oxide being reduced and is chosen to ensure that the carriermaterial is prevented from solidifying. For iron oxide ore reduction thetemperature is typically about 1300° C., which is high enough to promotereduction but not high enough to cause melting of the solids in thefloating charge layer.

The proportions of metal oxide to carbonaceous reductant depend againupon the metal oxide being reduced and also open the nature of thecarbonaceous reductant. However, the proportions employed will be suchas to given carbon monoxide. Thus, it will be usual to operate theprocess with stoichiometric excess of carbon relative to the amount ofoxide to be reduced.

In order to maximize the efficiency of heating of the molten carriermaterial in the heating zone it is necessary to prevent accumulation ofslag or solids in the heating zone. This can conveniently be effected byremoving the molten carrier from under the slag or any residual solidsin the charge arm before passing it to the heating zone.

The solid metallised raft produced in the reduction melt circulationloop described above next becomes the feed input for at least onefurther melt circulation loop in order to conduct continuous melting,slag separation and primary refining; the molten metal is then passed(eg. overflows or is siphoned) into at least one further meltcirculation loop to effect additional refining and compositionadjustment after which the molten metal continuously discharges eitherby overflowing or by siphonic removal so that it either becomes theproduct refined metal or alloy or optionally continues on for furtherrefining eg. in a packed tower for countercurrent contacting with argonor other suitable gas under reduced pressure to reduce impurities toultra-low levels.

The metallised raft comprised of solid metal, unreacted carbon, gangueoxide constituents possibly already reacted to form solid slag compoundsand solid fluxes to further aid liquid slag formation and metalrefining, is propelled out of the charge arm of the reduction loop intothe metallised raft melting and slag formation/separation arm of thefirst in-line refining loops, utilising the resultant drag force exertedon the raft by the molten carrier material in the charge arm of thereduction loop and possibly assisted by maintaining a shallow depth ofmolten carrier material on the cross-over interconnecting the two saidmelt circulation loops to ensure unimpeded continuous transference ofthe metallised raft onto the surface of a second molten carrier materialbeing circulated in a closed loop path serially through a topblown/flame impingement zone in order to melt the metallised raft, burnexcess carbon on one arm of the loop and a second arm on which heatingand additional refining processes such as desulphurisation anddecarburisation are conducted employing top blowing and liquid fluxcontacting in concert.

Separation of slag and metal occurs by the solid metal in the metallisedraft being assimilated into the molten carrier material and the slagforming a continuous molten layer floating on the molten carriermaterial down to the far end from the feed input where it eitheroverflows with the molten carrier material into a downstream zone whereslag can accumulate for intermittent tapping while the molten carriermaterial is removed from beneath the accumulated slag for continuouscirculation to the other arm of the melt circulation loop using agas-lift mechanism or siphon, whichever is the more appropriate.Alternatively, the overflow weir can be dispensed with and use made ofan electromagnetic dam as developed for continuous strip casting and thelike to keep the molten metal back whilst accumulating a pool of moltenslag which can be removed either continuously or intermittently bytapping.

Unlike the reduction loop, the molten carrier materials in the refiningloops are involved chemically in the process because the metal beingproduced is dissolved therein and impurities such as carbon, silicon,sulphur and phosphorus interact thermodynamically with each other inways that can influence the choice and effectiveness of refiningprocesses to which the molten carrier materials are subjected, it beingappreciated that because of the large melt circulation ratiosestablished in the refining loops and already referred to for thereduction loop, the molten product overflowing from these loops isvirtually the same as the molten carrier and the melt compositionthroughout a particular loop is effectively constant and temperaturedifferences in the bulk of the molten carrier throughout are relativelyminor.

An optional final step in the process for producing molten refined metalor alloy is to cause the molten metal, either overflowing or beingsiphoned out from the last of the melt circulation refining loops, toirrigate solid packing elements within a packed tower with molten metalflowing by gravity downwards in the form of droplets and rivulets with arising gas flow such that true countercurrent contact is maintainedbetween the gas and liquid phases with longitudinal or backmixingreduced to an absolute minimum. A preferred way of achieving this truecounterflow is to operate the tower under reduced pressure so that thedownward velocity of the liquid metal droplets and rivulets is less thanthe upwards velocity of the gas. For ultra-low carbon (ULC) steelproduction, the objective of this countercurrent contacting is to effectthe reaction between dissolved carbon and dissolved oxygen bymaintaining virtually zero concentration of gaseous reaction product,carbon monoxide, at the base of the solid packing where high purityargon is admitted.

For the UCL Tower Refiner any suitable relatively large packing elementsmay be used with 100-150 mm MgO or MgO/Al₂O₃ spinel being a preferredchoice. In considering the diameter of the ULC Tower it is important tokeep below the so-called flooding condition and also the gas phasepressure drop should only be a minor fraction of the total pressure. Forthe case of ULC steel, an elaborate vacuum system is not required andall that is needed is a water ring pump to exhaust the purge gas and itsassociated carbon monoxide at a pressure in the region of 100 mbar.

The method of this aspect of the present invention is suitable fordirect coal-based continuous steelmaking from fine coal and iron orefines. Although it has already been pointed out, the molten carriermaterial in the reduction loop is not formed directly from the chargematerial, it is convenient to use hot metal (hot molten iron, impure inthe as-smelted state) as the molten carrier material and to add orremove hot metal from the first loop as the on-going process demands.For the other loops, which constitute the continuous steelmakingprocess, however, the iron produced by reduction forms the carriermaterial, which is removed, preferably continuously, at a rate whichsubstantially balances the rate of iron production.

There have been many unsuccessful attempts in the past to developcontinuous steelmaking. Problems with refractories stand out as theprincipal contributors to unreliability and excessive operating costs inthose processes reaching semi-plant size trials. The source of theseproblems can be traced back to the aggressive behaviour of molten ironoxide towards refractories, exacerbated by reactor high intensity whenoxygen is used directly as in modern steelmaking. When oxygen is useddirectly to refine hot metal, subsurface formation of carbon monoxidebubbles generate mild explosions throwing molten slags and splashes ofmetal into the surrounding gas accompanied by further intense reactionof ejected droplets and copious generation of iron oxide fumes.

Besides attack on the refractories, the above scenario generatesaccretions and other problems which are incompatible with continuousprocessing. A more moderate approach is needed, preferably one whichtakes into account legitimate concerns about climate change andgreenhouse gas emissions as well as social issues relating to health andsafety, sustainability and the environment. Considerable benefits to allthese would stem from substitution of oxygen steelmaking with carbondioxide blowing for decarburisation coupled with efficient postcombustion to offset the endothermicity of the reaction chemistry.Oxygen should be used for post combustion of carbon monoxide generatedduring reduction, but first the carbon monoxide should be partiallycombusted with oxygen in a step-wise fashion to provide the gaseousoxidant for top blowing along with water vapour derived from hydrogen inthe reductant for steelmaking rather than using oxygen directly.Sequential combustion of the carbon monoxide also permits heat to begenerated for melting the metallised raft in the first of thesteelmaking loops, whilst maintaining iron as the thermodynamicallystable phase so iron oxide is not formed at unit activity and theproblem of refractory attack is greatly reduced. In addition, except foruncontaminated nitrogen from air separation, processes of the futuremust move ideally towards zero gas emission after sequestration of thecarbon dioxide. Clearly, oxygen processing is needed but not in the wayit is used in current steelmaking practice.

As we move towards a hydrogen economy, there will still be a need todecarburise iron melts containing minor amounts of carbon to producerefined steel. In this case the gas phase will be principally hydrogendiluted with nitrogen or just hydrogen. The gaseous mixture required forprimary and secondary steel making will be N₂/H₂O again ensuring that nofree oxygen comes into direct contact with carbon containing molten ironanywhere in the process.

It is now well established that interfacial chemical kinetics play animportant role in the reaction of carbon dioxide or water vapour withcarbon dissolved in molten iron during the decarburisation process ofprimary steelmaking. The slow chemical step involving dissociativeadsorption is not observed with oxygen, where gaseous diffusion andliquid phase mass transfer can both influence the rate ofdecarburisation. Accordingly, to moderate the steelmaking reaction,carbon dioxide and water vapour are the preferred oxidants. This isespecially the case when the molten iron being decarburised containsdissolved sulphur, as it undoubtedly will in any real steelmakingsituation. This is because sulphur slows down the dissociativeadsorption reaction on the gas-liquid interface.

The lesser intensity referred to above is turned into a positiveadvantage in the present invention, where it is imperative thatsub-surface nucleation and growth of carbon monoxide bubbles must not beallowed to occur in the interests of avoiding explosive ejection ofmetal droplets into the gas space followed by further intense reactionwith sparking and copious fume generation, all of which are incompatiblewith continuous processing as discussed earlier. This is a keyrequirement, which must be taken into account in specifying the methodsto be used in the melt circulation steelmaking loops for continuoussteelmaking without disruptive ebullition and ejection of melt frombubbles as they burst through the flowing melt surface withconsequential skull or accretion build-up on the walls and roof of thereactor, which would eventually necessitate shutdown of the continuoussteelmaking plant. In the present invention, this is secured by adaptingthe processing conditions so that the supply of carbon to the meltsurface by liquid phase mass transfer throughout all of the steelmakingloops is always adequate to prevent oxygen atoms from diffusing into themolten metal to such an extent that concentrations of both oxygen andcarbon in the bulk molten iron reach supersaturation levels sufficientto induce the decarburisation reaction to occur spontaneously beneaththe surface.

For the particular case of coal-based continuous steelmaking, twosteelmaking loops are preferred. The first is top blown to effectprimary decarburisation, whilst in the second loop what has been termed“open-channel” decarburisation is promoted under increased meltcirculation rate. For both loops, steady-state conditions areestablished such that gas phase mass transfer, interfacial chemicalkinetics and liquid phase mass transfer all balance each other.

Particular examples of the invention as applied to direct coal-basedcontinuous steelmalking will now be described with reference to theaccompanying drawings in which:—

FIG. 1 is a schematic general arrangement in plan view of the plant fordirect coal-based continuous steelmaking, when steel scrap, hotbriquetted iron (HBI), or direct reduced iron (DRI) are readilyavailable and their use is economical.

FIG. 2 is a schematic general arrangement in plan view for directcoal-based continuous steelmaking for a stand-alone plant based onvirgin ore as the only source of iron units.

Referring now to FIG. 1, the plant comprises six furnace hearths 1, 2,3, 4, 5 and 6, which are arranged in pairs to form three inter-linkedmelt circulation loops A (a charge reduction loop), B and C (first andsecond refining loops) formed by interconnecting the first and secondhearths 1 and 2 (constituting a charge reduction and adesulphurisation/heating zone respectively), the third and fourthhearths 3 and 4 (constituting a melting zone and adesulphurisation/decarburisation zone) and the fifth and sixth hearths 5and 6 respectively. Under steady operating conditions, molten metal iscaused to overflow or be otherwise taken out of the second and thirdloops B and C by conductively heated siphons 7 and 8 so that moltenmetal issuing from these is equivalent to the metal in the compositecharge initially added to the top surface of the molten carrier materialat the upstream end of the first hearth 1, together with any scrap orpre-reduced material added to the circuit and shown in FIG. 1 as 12. Atthe downstream end of the first hearth 1, a channel or ramp 8 isprovided to permit solid material (metallised raft) floating on thesurface of the molten carrier material in the first hearth 1 to bepropelled or projected into the third hearth 3 onto the surface of themolten carrier material in the first refining loop B, along with alesser amount of molten carrier material from the reduction loop A,corresponding to new metal assimilated into the carrier material inreduction loop A, resulting from any smelting reduction taking placebetween the floating charge layer, ultimately becoming an agglomeratedsolid structure referred to as the metallised raft as it progressesdownstream in the first hearth 1.

Forced melt circulation in all loops is effected by R-H type snorkels 9connected to vertical bodies 10 linked to each other by horizontalmembers 11 to form vacuum-tight refractory-lined vessels, which canfunction either as gas-lift pumps or siphons depending on whether or notan inert gas is injected into the upleg snorkels. Under reduced pressuremelt is drawn up into both snorkels 9 and the lower regions of thevertical bodies 10 and the horizontal members 11 in each unit to form achannel through which the melt traverses as it flows from one hearth toanother along the horizontal member 11. These vessels can be eitherlowered so that the snorkels 9 are immersed in the molten carriermaterials or raised for stand-by or replacement with preheated units ona scheduled maintenance basis.

The very much smaller siphons 7 and 8 have similar features to the unitsdescribed above, but because the melt flow rates in these siphonscorrespond to the actual metal production rate, steps may need to betaken to independently heat the flowing metal by direct resistanceheating or so-called conductive heating.

The composite charge 12, comprised of well-mixed iron ore fines, finecoal and preferably burnt lime flux, is distributed uniformly onto thesurface of the molten carrier material towards the upstream end of thefirst hearth 1 to form a floating charge layer 5 to 10 cm in thickness,whilst being heated from beneath by the molten carrier material. Coalvolatiles and reaction product gases are discharged from the gas spaceabove the floating solid charge into a refractory-lined gas header duct13, which forms the manifold for an array of top blowing lances withconcentric controlled oxygen admission so that the metallised raftfloating in the third hearth 3 is melted to form a liquid slag layer,whilst the metallics are assimilated into the molten carrier material.Slag is removed at the downstream end of the third hearth 14 eithercontinuously or intermittently from a pool of molten slag formed whenthe molten metal is held back by an electromagnetic device or damsimilar to that being developed for continuous casting applications.

The gases flow from the third hearth 3 into a hot gas clean-up (HGCU)facility, which incorporates a combined liquid-metal quench anddesulphurisation tower 15, a turbo-booster 16 and a liquid-metal basedgas reheater 17. In other embodiments (not shown), the gases arearranged to flow into the HGCU facility from the first hearth.

With its pressure now increased, the hot gas from the HCGU flows into arefractory-lined header duct 18, which forms the manifold for an arrayof top blowing lances with further controlled concentric oxygen additionso that carbon dioxide and water vapour become the oxidant gases ratherthan oxygen itself for primary or major decarburisation of the moltencarrier metal in the fourth hearth 4 using direct flame impingement toprovide both the thermal requirements and the gaseous reactant for theendothermic decarburisation reaction.

Besides providing the principal decaiburisation requirements, the fourthhearth 4 is used to effect flux-based desulphurisation, whilst thedissolved oxygen content of the iron is relatively low and the carbonlevel sufficiently high to sustain this requirement. Accordingly,desulphurisation flux is added at the upstream end of the fourth hearth19 and removed at its downstream end 20.

The hot gases from the fourth hearth 4 discharge into a refractory-linedheader duct, which forms the manifold for an array of top blowing lances21 with further concentric controlled oxygen addition to effect meltingof preheated steel scrap or prereduced material in a pool of moltenmetal 99, which overflows liquid scrap into the fifth hearth 5 where themelt is flowing at an accelerated rate in the end in which open-channeldecarburisation of the melt siphoned from the fourth hearth 4 iseffected by contacting the melt in the fifth hearth 5 with the oxidisinggases leaving the pool melter 22 under turbulent flow conditions.Because of the lesser amount of decarburisation that occurs in thesecond refining loop C, the sensible heat of the gases leaving the poolmelter 22 is more than adequate to provide the endothermic requirementsof secondary decarburisation before the gas flows into the sixth hearth6 via the refractory-lined cross-over gas duct 23. Because the carbonlevel throughout the second refining loop C is low, the melt iseffectively steel for general purpose applications and the dissolvedoxygen level is high enough to conduct flux-based dephosphorisation byadding powdered flux at the upstream end 24 of the sixth hearth andremoving flux at the downstream end 25.

After passing along the length of the sixth hearth 6, the hot gasesdischarge into a refractory-lined transfer duct 26, which becomes themanifold 27 for an array of top blowing lances with major concentricoxygen addition (or other suitable configuration such as mutuallyopposed jets) to effect combustion to completion of the gases emittedinitially from the reduction loop A via the off-take 13 into the firstof the refining (steelmaking) loops B. This major combustion ultilisesdirect flame impingement onto the molten carrier material in the secondhearth 2 to contribute towards the post combustion energy needed tosustain ironmaking in the first hearth 1.

A small amount of a desulphurisation flux, which wets the melt surfaceand spreads across it to form a continuous thin film on the surface ofthe molten carrier material, is added at the upstream end 28 of thesecond hearth 2 and taken away at its downstream end 29. Thiscontributes to desulphurisation in the circuit, but most importantlyraises the emissivity of the molten carrier material and so considerablyenhances heat transfer by radiation. The very hot combustion gases aredischarged into a refractory-lined off-take 30, possibly with stavecooling within the refractory lining, directly into a countercurrentradiant scrap preheater 31, in which a sloping hearth plus a mechanicaldevice is used to contact selected steel scrap 39 to effect major scrappreheating and possible partial melting, before the scrap or otherreduced material passes into the pool melter 22. The off as from thescrap preheater leaves the plant at 32 and can be used in a heatrecovery steam generator (HRSG) for advanced power generation.

Vacuum degassing of steel using the RH process is established practicethroughout the world. By injecting argon into a single upleg, moltensteel circulation at rates up to about 200 tonne per minute can beobtained and this is now regarded as state of the art in the steelindustry. However, elaborate vacuum plant as used in RH degassing is notneeded for the melt circulation systems in the present invention. Argoncan be used as the lift gas but it may be preferable to use thedesulphurised gas arising from ironmaking whilst it is still notoxidising iron, because argon is needed elsewhere in the circuit,particularly if ultra-low carbon steel (ULC) is to be producedcontinuously using the molten steel discharging from 8 as the feed intothe Tower Refiner described previously.

It must be stressed that the overall energy implications of meltcirculations as shown schematically in FIG. 1 are minimal. To circulatemelt thorough a closed loop path is a function of the product of therate at which melt is circulated times the total liquid head to bepumped against. For the present case, differences in level throughoutthe whole system are mainly due to frictional effects and otherphenomena associated with open channel fluid flow at high throughput.Accordingly, the pumping heads can be kept small by design so that verylarge circulation rates can be employed for the various molten carriermaterials without consuming undue amounts of energy.

For the steel industry world-wide, the ultimate vision is perpetualrecyclability with steel scrap being recycled again and again with inputfrom iron ore only utilized to accommodate growth in demand. Thisembodies the desirable concept of environmental sustainability,requiring conservation of natural resources (iron ore and coal),minimizing the production of CO₂ in the first instance and sequestrationof whatever CO₂ is produced to combat climate change and global warming.

To maximize scrap melting or charging of DRI, HBI etc. if so desired,all of the gases emitted from the ironmaking loop A need to be utilizedtogether in both of the steelmaking loops B and C. Ideally, the smeltingreduction gases in total need to be passed directly to the metallisedraft melting arm (hearth 3) of the primary steelmaking loop A andpartially combusted with oxygen with direct flame impingement using anarray of top blow lances to effect iron melting and liquid slagformation as the metallised raft floats along with the circulating ironcarbon melt with phase disengagement completed when a molten slag layerfree from of associated iron is established towards the downstream endof the arm in question, in advance of slag removal from the circuit.

To ensure high iron recovery, metallic iron is maintained as thethermodynamically stable phase by controlling the oxygen addition inboth the smelting reduction arm (hearth 1) of the ironmaking loop A andthe metallised raft melting arm (hearth 3) of the primary steelmakingloop B.

The principal supply of heat for smelting reduction is by meltcirculation in the ironmaking loop A. Soot formation as coal isdevolatilised and reduction commenced is precluded by minor oxygenadmission above the charge material as it floats down the charge arm(hearth 1) of the ironmaking loop A but the mixed gases are neverallowed to become oxidizing to metallic iron, either here or after theflame impingement top blow arrangement for melting the metallised raftin the first steelmaking loop B. This is a unique feature, only feasiblein a melt circulation system, and guarantees very high iron recoverywithout oxidation losses. It assumes, of course, that the retention timein the ironmaking arm (hearth 1) is adequate to achieve a high degree ofmetallisation in the first instance. Once solid metallic iron is formedit will be impossible to lose any of this by solution in the slag, norcan it be oxidized within the ironmaking furnace itself.

Direct flame impingement was selected so that metallic iron at unitactivity can be melted directly without relying on dissolution in theFe—C melt. Similarly, the processes involved in slag formation areaccelerated by this approach.

Very little of the heat required in the steelmaking loops is derivedfrom the sensible heat of the circulating melt. This has importantimplications relating to the design of the above bath enclosures. Lowthermal mass insulating materials, commercially available, areincorporated throughout the walls and roofs of both steelmaking loops Aand B and for the charge side (hearth 1) of the ironmaking loop A. It isonly the desulphurisation/heating arm (hearth 2) of the ironmaking loopA which may require a high specification refractory enclosure, althougheven here it may be feasible to use low thermal mass and relativelyinexpensive insulating materials currently being developed. Obviouslythis must have a major impact in terms of capital cost savings. It alsofacilitates easy access to the hearths, if needed, and generally reducesthe structural requirements because the materials involved arelightweight in comparison with brickwork or castable refractories.

The hot gases undergo hot gas cleanup (HGCU) involving a liquid metalquench, which incorporates sulphur removal to a very high level as wellas removal of particulate solids, before being boosted in pressure toaround 0.5 to 1 bar gauge and reheating again by liquid metal directheat exchange. The gases are returned to the steelmaking loop B at about1350° C. for partial combustion with oxygen to provide the chemical andthermal requirements for major decarburisation of Fe—C melt by directflame impingement under non-splashing top blow conditions on thedesulphurisation flux arm (hearth 3) of the primary steelmaking loop B.The momentum of the jets clears the flux away and mixes it to assistdesulphurisation, whilst promoting liquid phase mass transfer ofdissolved carbon from the bulk of the liquid metal to the interface,where the endothermic reactions with CO₂ and H₂O tale place.

The top blow lances are designed so that the supply of gaseous oxidantto the liquid metal interface at steady state is balanced by the supplyof carbon by liquid phase mass transfer. Allowance is also made for thekinetics of the dissociative adsorption of both CO₂ and H₂O, includingthe inhibiting effects of sulphur. The resulting steady state conditionsare such that sub-surface CO bubble formation cannot take place.Laboratory experiments using the electromagnetic levitation techniquehave established that sparking or copious fume generation will not occurunder these relatively mild conditions.

The gases during decarburisation may gain enough CO by reaction thatthey may leave the primary steelmaking loop A with metallic iron againas the stable thermodynamic phase. Further controlled addition of oxygenis made to these gases in another top-blow flame impingementarrangement, this time directed at a pool of molten iron, which isessentially a cul-de-sac off the second refining loop-C for what istermed open-channel decarburisation. This pool receives partially meltedsteel scrap from a radiant heater in the form of a sloping hearth,leading directly into the pool and fired by the very hot gases from thesecond hearth. The overflow from the pool is liquid scrap, which joinsthe main melt circulation flow, any non-metallic residues beingincorporated into the dephosphorisation flux layer as it floats downthis arm (hearth 5) of the open-channel decarburisation loop C. The hotgas mixture then transverses the length of the open-channeldecarburisation arm (hearth 5), where a clean metal surface permitsfurther decarburisation to a low level without sparking or fume emissionand then the hot gases are ducted to the heating arm 2 of the ironmakingloop A where further oxygen is added for full combustion before exitingto the sloping hearth radiant heater for partial melting of steel scrapat a temperature at between 1700-1800° C. at the inlet withcountercurrent contacting of the incoming scrap feed

The gases ultimately leave the radiant heater at around 850° C. fortransmission to the heat recovery steam generator (HRSG), which togetherwith the steam tubes associated with the various frozen shellsconstitute the steam boiler for advanced power-generation.

The embodiment described above when applied to plant for continuouslyproducing steel from iron ore fines, when steel scrap, HBI or DRI areboth readily available and economic to use, would typically producethree tonnes of steel product from every one tonne of virgin iron unitsincorporated into the composite mixed feed.

There will always, of course, be a need for creation of value-addedproducts by primary production away from centres of high population andscrap availability and, ideally, new technology must be able to switcheasily from high scrap utilisation to 100 pct virgin raw materials, ifthe situation so demands. It is this in-built flexibility that is a keyfeature of the present invention. In place of the scrap meltingarrangement shown in FIG. 1, the gases in 21 flow directly to the fifthhearth 5, the pool melter 22 and the radiant scrap preheater are botheliminated.

This means that extra combustion is available in the second hearth andto prevent this from over heating it is necessary not to fully combustthe gases therein but to pass the very hot gases now containing residualCO and H₂ directly to a Waste Heat Boiler (WHB), probably with CO₂recirculation to moderate temperatures. In the event, more electricitythan needed for air separation and CO₂ liquefaction will be generated.Depending on the proximity of other plant nearby or other in-plantusage, some of the power generated would be available for export to thegrid if such an opportunity exists in centres away from denselypopulated areas. Alternatively, if steel scrap is not readily availableand the use of HBI or DRI is uneconomic, then this becomes the preferredoption with export of electricity to the national grid.

Referring now to FIG. 2, a stand-alone plant retains the principalfeatures of having six furnace hearths, which are arranged in pairs toform three inter-linked melt circulation loops. Except for the changesdiscussed above in relation to the non-availability of steel scrap atthe right price and assuming that using HBI or DRI is also not economic,the other features in FIG. 2 are essentially the same as those shown inFIG. 1 and corresponding reference numerals are used accordingly. Forthe stand-alone plant, however, since excess thermal energy isavailable, it may be desirable to incorporate calcination of limestone33 to burnt lime 34 into the circuit. This is shown schematically inFIG. 2 as an adjunct downstream of the off-take 30 of the very hot gasesfrom the post combustion arm (hearth 2) of the ironmaking loop A. Itinvolves recycling a stream 35 of CO₂ and H₂O at a temperature at around125° C. from the exit flue of the WEB to attemperate the gas leavinghearth 2 through the off-take 30, which is split into two streams, intoone of which powdered limestone 33 is injected. Sufficient residencetime is provided in the gas flow path within the calciner 36 to produceburnt lime 34, which is recovered by a cyclone separator 37 and thecombined gases 38 then proceed to the WEB. This burnt lime is one of theprincipal components of the mixed composite charge 12 and may also beused in preparation of the desulphurisation flux 19 and thedephosphorisation flux 24.

The method according to the present invention is not solely applicableto iron oxide ores such as haematite, but is also applicable forexample, to the production of refined metal or alloy from the following:

(1) Chrome ores to form ferrochromium alloys.

(2) Chromite together with nickel sulphide to make master alloys forstainless steelmaking.

(3) Chromite and nickel oxide for direct stainless steelmaking.

(4) Nickel laterites to form ferronickel directly.

(5) Deepsea manganese nodules to form copper-nickel-iron alloy (furnacealloy).

(6) Manganese ore to ferromanganese alloys.

1. A method of reducing metal oxide material to metal comprising thesteps of:— (i) circulating a first molten carrier material in a closedsolid-state reduction loop serially through a charge reduction zone anda combined melt desulphurisation and heating zone; (ii) circulating asecond molten carrier material in a closed melt refining loop seriallythrough a melting zone and a desulphurisation/decarburisation zone (iii)introducing a mixed composite charge comprising a metal oxide and acarbonaceous reductant onto the surface of the carrier material in thereduction zone of the solid-state reduction loop, (iv) reducing saidmetal oxide to solid metal by the carbonaceous reductant in thereduction zone, the metal oxide and carbonaceous reductant beingintroduced in step (iii) in proportions such that the carbon from thecarbonaceous reductant is converted to carbon monoxide, hydrogen gasalso being produced; (v) separating a metallised raft containing thesolid metal from said molten carrier material by projecting saidmetallised raft into the refining loop upstream of the heating zone ofthe solid-state reduction loop so that the surface of the first moltencarrier material which is circulated to the heating zone issubstantially free of solid material; (vi) mixing the carbon monoxideand hydrogen formed in step (iv) and partially combusting the mixture inthe melt refining loop in order to satisfy the thermal requirements formelting the metallised raft and for the chemical reactions taking placein the desulphurisation/decarburisation zone and providing the gaseousoxygen mixture containing no free oxygen in order to effectdecarburisation; (vii) optionally carrying out further partialcombustion of the carbon monoxide and hydrogen to melt solid scrap ordirect reduced material; (viii) after steps (vi) and (vii), mixing thecarbon monoxide and hydrogen with oxygen and completely combusting themixture well above the melt in the heating zone of the solid-statereduction loop so that although direct flame impingement is facilitated,no unreacted oxygen comes into contact with the melt in providing thefull thermal requirements for reduction of the oxide charge materials inthe solid state by transfer to the first molten carrier material whichis recirculated to the solid-state reduction zone; (ix) submitting thecarbon monoxide and hydrogen to hot gas cleanup for removal of sulphurand solid particulates prior to step (vi), (vii) or (viii); (x) meltingthe metallised raft in the melting zone of the refining loop andrefining the molten metal in the desulphurisation/decarburisation zone,and (xi) removing the metal so refined from the refining loop.
 2. Themethod as claimed in claim 1, wherein the composite charge additionallycomprises a flux.
 3. The method as claimed in claim 2, wherein said fluxis selected from lime, limestone and dolomite.
 4. The method as claimedin any preceding claim, wherein said metal oxide is selected from one ormore of iron, chromium, nickel and manganese oxides.
 5. The method asclaimed in any preceding claim, wherein the composite charge added instep (iii) forms a layer having a thickness of from 5 cm to 20 cm,preferably 5 cm to 10 cm, on the first carrier material.
 6. The methodas claimed in any preceding claim, wherein the composite charge is atleast partially compacted prior to step (iii).
 7. The method as claimedin any preceding claim, wherein the ratio of carrier material to metalto be produced is from 100:1 to 500:1, preferably 200:1 to 400:1.
 8. Themethod as claimed in any preceding claim, wherein the charge reductionzone of the solid-state reduction loop is constituted by a first hearthand the desulphurisation/heating zone is constituted by a second hearth,a flowpath being provided between a downstream end of the first hearthand an upstream end of the second hearth and between a downstream end ofthe second hearth and an upstream end of the first hearth.
 9. The methodas claimed in any preceding claim, wherein the melting zone of therefining loop is constituted by a third hearth and thedesulphurisation/decarburisation zone is constituted by a fourth hearth,a flowpath being provided between a downstream end of the third hearthand an upstream end of the third hearth and between a downstream end ofthe fourth hearth and an upstream end of the third hearth.
 10. Themethod as claimed in claim 9, wherein said flowpaths are provided by gaslift pumps or siphons.
 11. The method as claimed in any preceding claim,wherein the velocity of the carbon monoxide produced in step (iv)increases toward the downstream end of the charge reduction zone. 12.The method as claimed in any preceding claim, wherein hot gas clean upis effected between steps (iv) and (viii).
 13. The method as claimed inany preceding claim wherein any slag produced in step (iv) is removedfrom the carrier material before the latter is passed into thedesulphurisation/heating zone.
 14. The method as claimed in anypreceding claim, wherein step (v) is achieved by maintaining a shallowdepth of the first molten carrier material in a crossover between themelt circulation loop and the refining loop, so that the metallised raftfloats on the first molten carrier material and is transferred to therefining loop by drag forces exerted by the first molten carriermaterial.
 15. The method as claimed in any preceding claim, wherein themetal removed from the refining loop in step (viii) is passed to anadditional refining loop for further refining.
 17. A metal refiningapparatus comprising:— (i) a solid-state reduction loop comprising firstand second hearths, a continuous fioVpath CXIsLing between a downstreamend of the first hearth and an upstream end of the second hearth and adownstream end of the second hearth and an upstream end of the firsthearth, (ii) a refining loop comprising third and fourth hearths, acontinuous flowpath existing between a downstream end of the thirdhearth and an upstream end of the fourth hearth and a downstream end ofthe third hearth and an upstream end of the fourth hearth, (iii) meansfor introducing a composite charge into the upstream end of the firsthearth, (iv) means for transferring, in use a metallised raft floatingon a molten carrier material from the downstream end of the first hearthto the upstream end of the third hearth, (v) ducting for transportinggases generated in the first hearth to the third hearth for partialcombustion, (vi) ducting for transporting the partially combusted gasesin the third hearth to the fourth hearth for further combustion, (vii)ducting for transporting the partially combusted gases in the fourthhearth to the second hearth for complete combustion with oxygen, and(viii) a hot gas clean up unit in the gas flowpath between the first andthird hearths or between the third and fourth hearths.